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By Edward Martinez

The reactions occurring in the copper segregation process were studied by heating mixtures of chryso-colla, salt, and a reducing agent. The techniques used in this investigation were differential thermal analysis (DTA) and thermal gravimetric analysis (TGA). The initial chloridizing of the oxide copper occurs below 500°C and cupric chloride may be the compound formed. Volatilization and reduction of the copper chloride take place above 550°C, although these reactions do occur to some extent below 550°C. A difference in the distribution of the metallic copper in the reducing agent was found when sodium chloride was used compared to magnesium or calcium chloride. Segregation was obtained with mixtures of cupric oxide, salt, and charcoal with no quartz or silicate present. Preliminary tests indicate that segregation may occur even in the absence of water vapor. The DTA and TGA techniques were found to be of great value in studying these reactions. The copper segregation process is a method for treating oxide copper ores. The ore is heated, most commonly in a kiln, with a reducing agent and a halide salt, generally sodium chloride, to approximately 700°-750°C and held for 30 to 60 min. The addition of the salt results in the migration and reduction of the oxide copper to metallic copper on the surface of the reducing agent rather than in the original mineral structure. The metallic copper can then be recovered by a subsequent leaching or flotation step.1-4 Experimental work in 1923 on oxide copper ores from the Sagasca mines in Chile led to the discovery of the reduction process. It was found that during the reduction of the copper mineral with coal, at about 700°C, the copper migrated to the surface of the coal instead of remaining disseminated in the ore. The unusual behavior of the copper, which caused it to migrate from the ore particles to the coal, was traced to the presence of a small quantity of sodium chloride occurring in the ore. Further experiments with oxide copper ores proved that the presence of a halide, either occurring naturally or admixed with the ore, caused the copper to migrate from the ore particles and to segregate in the charge. The ratio of salt to reducing agent may vary somewhat, but in general 0.5% to 1.5% sodium chloride and 0.5% to 1.0% coal or coke have been mixed with the ore. Less than stoichiometric amounts of salt are required to convert all the oxide copper to a chloride, so that a cycle of reactions appears to be occurring. Several investigations reported in the literature have proposed reactions to explain the mechanism of the process.5-10 There appears to be general agreement that water is necessary for the reactions to occur; that once the sodium chloride reacts with silica or silicates in the ore, the hydrochloric acid formed will convert the cupric oxide in the mineral structure to cuprous chloride. However, there appears to be a conflict of opinion on the reactions involved in the reduction of the cuprous chloride to metallic copper on the surface of the carbon. Rey has indicated that salt, water vapor, and

Jan 1, 1968

By J. Steininger

The Hallimond tube flotation of sphalerite and pyrite with sulfhydryl collectors has been found to be inhibited near neutral pH in the presence of copper salts. Alkalimetric titration studies of the reactions between collector and activator indicate that this effect is apparently related to the formation of basic copper collector complexes. The conditions of formation of these compounds and their stability are in good agreement with the depression observed in flotation experiments. The general procedure employed in the flotation of zinc sulfide minerals requires activation of the mineral surface by heavy metal ions, usually copper, which react more easily than zinc with the common short-chain ionic sulfhydryl collectors.' Commercially, zinc flotation is carried out under alkaline conditions (pH 7 to 10.5), but small-scale experiments indicate that zinc sulfide minerals can also be activated and collected under acid conditions. Selectivity, however, becomes poor at low pH and undesirable materials such as pyrite, silicates and carbonates tend to report in the float. Several years ago, Wark and Sutherland~, measured the contact angles of air bubbles on copper-activated sphalerite immersed in various sulfhydryl collector solutions. When represented, as in Fig. 1, on pH-sodium cyanide diagrams, the results are characterized by 'islands' of non-contact near neutral pH and

Jan 1, 1969

By James E. Lawver, James L. Wright

This paper discusses the design and calibration of a Faraday pail for measuring electric charge. The paper also shows that at least for two minerals, quartz and calcite, the phenomenon that Johnson termed "reversibility" can be explained in terms of the contact electrification between the minerals and the metal rotor of an electrostatic machine. Each year the free world produces about 13 million tons of mineral concentrates by electrical separation. However, despite the rather general acceptance of electrical concentration as an industrial process, research in this area has lagged far behind the efforts devoted to competitive processes such as froth flotation. The purpose of this paper is to describe the design and calibration of a Faraday ice pail for measuring charge density on mineral grains and to show some of its applications in the area of electrostatic and electrodynamic research. THE DESIGN AND CALIBRATION OF FARADAY PAILS FOR ABSOLUTE CHARGE MEASUREMENT The electric charge on mineral grains at a given time during an electrical concentration process is de- pendent on many factors and cannot, in general, be accurately calculated a priori. It is thus important to devise a method of measuring absolute charge. One method of measuring electric charge is to note the potential rise of an electrometer when charged particles drop into a Faraday pail. Two types of Faraday pail have been constructed and calibrated. The first type, Faraday Pail I (Fig. 1), is a vertical pail originally designed to measure the charge due to contact electrification between mineral grains and a metal rotor or due to conductive induction when mineral grains are passed over a rotating cylinder in the presence of an electric field. The second type, labeled Faraday Pail II (Fig. 2), is a horizontal pail designed to determine the charge distribution of mineral grains after they pass through the separating zone of the electrical concentrator shown in Fig. 4. The main advantage of the horizontal pail is its low head-room requirement and the long opening which enables the experimenter to sample the entire width of the stream of minerals being separated. The Faraday pails were calibrated in the following manner: A calibration head was constructed in the form of a hollow brass sphere that could be held in an insulated metal cylinder by vacuum (Fig. 1). The capacitance of the sphere was calculated by measuring the capacitance of the sphere and cylinder combined and that of the cylinder alone and subtracting the latter measurement from the former. The sphere was then charged to a known potential (V) and allowed to drop into the Faraday pail which was attached to a Keithley Model 610A electrometer. This instrument has an accuracy of 2% full scale and has

Jan 1, 1969

By K. E. Merklin, F. D. DeVaney

In the past most of the research in pelletizing has as its aim the production of a physically strong pellet. Now that this problem is solved, more attention is being paid to uniformity in grade and chemical composition. Consideration has been given in recent years to a study of additives which will make pellets partially or fully fluxed. Blast furnace tests show fluxed pellets to have advantages over both fluxed sinter and acid pellets. Lower slag volumes are used more frequently in the blast furnace and the uniformity of the charge, both physically and chemically, is becoming increasingly important. A considerable improvement can be secured by having all of the flux uniformly distributed through the pellets in a calcined state. A fluxed pellet, therefore, should be a superior product. The use of an ore burden that contains an appropriate quantity of fluxing materials is not new to the blast furnace industry. For years the Minette ores of France and Germany, the Clinton ores of the Birmingham, Ala., areas, as well as others, which are low in iron but contain calcium carbonate in sufficient quantities to give them self-fluxing properties, have proven economical. The adaption of the sintering process to the preparation of ore burdens provided a means for artifically associating the fluxstone with the iron ore prior to charging to the blast furnace. In 1944 Torpl discussed the effects of using lime in sinter made in Sweden on the sinter's metallurgical performance in a blast furnace. Hessle2 reported in 1945 on the advantages of using lime in sinter to increase the sintering plants' capacity and reduce fuel requirements. Robinson3 discussed the use of lime in Swedish sintering practice at the 1950 Blast Furnace, Coke Oven, and Raw Materials meeting. One of the suggested advantages was the combining of lime with silica to prevent the formation of a fayalite slag, which is difficult to reduce in the blast furnace. Nilson4 at the same meeting commented briefly on experimental work that had just started in the U. S. on the use of lime in sinter. The literature tends to indicate that most of this early work was directed towards improvement in sinter physical quality, and the fact that the feed to the blast furnace was partially fluxed was incidental. It was not long, however, before blast furnace advantages were being claimed. The matter was forcibly called to the attention of North American operators by reports of remarkable production records established by the Russians using fluxed sinter. The leading step was taken on this continent by the Steel Company of Canada in 1957; McMahon5 reported on the marked improvement in performance of an 18-ft blast furnace using fluxed sinter at its Hamilton works. Other North American operators quickly followed theStelco lead.

Jan 1, 1963

By Vernon F. Swanson

An equation has been developed which will permit the direct determination of free settling velocity of any sized particle encountered in mineral benefici-ation. The equation is based on Newton's Law with the friction factor defined as a function of the laminar boundary layer. Development of the equation is described first from an empirical trial-and-error approach. Examples are given of actual settling data from the literature versus calculated data to prove the validity of the new equation. Newton's law of resistance of a fluid to flow around a sphere is expressed in Eq. 1. Free settling velocities can be determined by setting Eq. 1 equal to the gravitational force on a particle and solving for V. Experience has shown that when the solids are greater than 1 mm in diam, or when Reynold's number for the system is greater than 800-1000, the friction factor term, fD, essentially becomes a constant, and Eq. 2-b becomes Newton's equation for free-settling and is applicable under the conditions stated. Another special condition exists for particles smaller than 0.05 mm or when Reynold's number for the system is less than three. In this case fD=24/NRor24µ/pL,dV,. Eq. 2-b can then be rewritten as

Jan 1, 1968

By H. Evans, L. A. Hunt

The raw materials used in the blast furnaces at the Geneva Works of U.S. Steel Corp. have a high degree of variability in size-consist and chemical content. To overcome the problems caused by the use of highly variable raw materials in the blast-furnace operation, large blending piles are used. An investigation of the blending system showed that variability was significantly reduced by the use of these blending piles. This paper describes the blending system, the method used to evaluate the effect of the blending system, and the results of the investigation. The remarkable improvement in blast-furnace performance since World War II has been in large measure due to the use of beneficiated burdens to reduce the variation of both chemical analysis and size consist. This practice allows the operator better control of his blast furnace, enabling him to operate more economically and to produce chemically uniform hot metal. Planning for uniformity usually begins during mining by extracting ore from several locations within a pit to produce a final product with a specified analysis. Another approach is to stockpile ores having widely varying analyses in individual piles and then drawing the required amounts from each stockpile to provide a blend with a predetermined analysis. A further improvement in blending procedures is to take ores from several sources, bedding them horizontally in layers and reclaiming by vertical slicing from the blended pile. The last procedure, when properly executed, is one of the best methods for achieving maximum uniformity with highly variable ores. U.S. Steel Corp.'s Geneva Works initially obtained iron ore for their blast furnaces from the Cedar City area of southern Utah. These ores were formed primarily by the replacement of limestone and siltstone formations with solutions containing iron and several other elements. The original iron content of the ores in the ground varied from a high of 70% to practically zero with a minimum acceptable iron content for shipping of 37%. Substantial amounts of silica, calcium, magnesium and alumina were found in the ores along with minor amounts of copper, lead, zinc, nickel, titanium, manganese, phosphorus and sulfur which were sometimes present in detrimental amounts. Iron-to-silica ratios varied considerably throughout

Jan 1, 1968

By Gordon E. Agar, David M. Hopstock

This paper describes the procedures and results of a series of experiments conducted to determine the depressing effect of sodium, calcium and magnesium ions on the amine flotation of quartz. It also presents an analysis of the experimental results by the electrical double layer theory. The flotation of quartz with primary amines has been practiced on a large scale in the phosphate industry for many years. Some glass sand —and recently iron ore — has also been treated by this flotation technique. These applications have been highly successful and demonstrate that a complete mastery of all the details of an operation is not needed to achieve an economic success. There are times, though, when it becomes important to recognize, both qualitatively and quantitatively, the effects of some previously ignored variables. The effects of cations on the amine flotation process, particularly those ions that contribute to water hardness and abound in sea water, are a case in point. In conjunction with the development of a cationic flotation reagent for separating quartz from apatite, it became necessary to quantify the influence of calcium and magnesium on the separation. In order to strengthen and broaden current understanding of the flotation process, the results were interpreted by means of the electrical double layer theory. THEORETICAL BACKGROUND Assume that, given the same equipment, procedure, and time, the recovery of quartz by amine flotation is a function of only the adsorption density of aminium ions in the Stern layer, i.e., that for any given adsorption density there exists only one recovery percentage. By the Stern-Grahame treatment of the electrical double layer,'-' for any ionic species in a solution in equilibrium with quartz,

Jan 1, 1969

By J. A. Herbst, G. E. Agar

The effect of fluid viscosity on the classification of solids in a liquid-solid cyclone was investigated. The separation size was found to be proportional to Additionally, it was found that the prediction of flow rate through a cyclone is dependent on a Reynolds criterion. Above a critical value, the flow can be predicted by a simple equation, whereas, below this value a variable orifice coefficient must be introduced. INTRODUCTION The solid-liquid cyclone is widely used as a device for classification, thickening, and heavy media separations. The more important variables and how they affect the performance of a cyclone are known and have been the subject of several investigations.' In general, these studies have paid little attention to the viscosity and those which have, introduced it through Stokes settling considerations. Systems which contain appreciable amounts of swelling clays exhibit viscosities much greater than water, and in these systems the viscosity term becomes important in determining the separation size. The work described here was done to determine in a Newtonian system the dependence of separation size on viscosity. BACKGROUND The separation size for a cyclone classifier has been related to the parameters of the cyclone by an equation of the form: where a is probably between 1.3 and 1.5, ß is between 0.50 and 0.57, and 6 has a theoretical value of 0.5. where y has a value between 1.8 and 2.0. Although these relationships are empirical, equations that differ only in the exponents have been derived from theoretical considerations. Those interested in the origin of these equations are referred to some of the original work.6,7 MATERIALS AND PROCEDURE The laboratory apparatus used in this study is pictured diagramatically in Fig. 1. Feed slurry was pumped from an agitated 15-gallon plastic tank by a centrifugal pump driven by 3/4 hp motor. The static pressure of the feedstream and thus the pressure drop through the cyclone was measured on a gage located as close to inlet of the cyclone as possible. When samples were not being taken, overflow and

Jan 1, 1967

By R. C. Ferguson, Harlowe Hardinge

Laboratory and plant data covering 12 different operations show that lower than "standard" ball mill speeds increase grinding efficiency. In the case of high pulp-level mills, the gain is so great that the increase in capital cost of the larger lower speed mill will pay for itself in less than a year's time. THE object of this paper is to show the economic advantage of operating ball mills at relatively slow speeds. Although many operators know that grinding costs are reduced by operating mills at low speeds, they seldom do so for a variety of reasons. When capital is limited, first cost is of primary importance, and installation of the less costly, highspeed mill is favored. When materials are in short supply, as during a war period, capacity rather than economy is the primary consideration. Existing mills are then speeded up to provide the desired increase in capacity, and new mills are purchased without consideration of maximum grinding efficiency. When the mill speed is increased to secure more production, the increase in grinding costs is often obscured by the total benefit derived from the additional output. This increase in grinding cost is seldom recognized as being sufficient to be worth correcting. For these reasons one may be lead to the erroneous conclusion that slow-speed operation is only occasionally applicable. The effect of mill speed on grinding efficiency has been studied in laboratory tests."! The results of some of these tests are given in tables I and I1 and fig. 1. In these tabulations, batch tests using ore charges of 125 to 200 lb were considered comparable to high pulp-level operation and charges of 50 to 100 lb comparable to low pulp-level operation. For the high pulp-level operation, these tests indicate the existence of a maximum in efficiency for operation at a speed of 50 pct of critical, both for the soft (dolomite) and hard (chert) ore. In low pulp-level operation, a decrease in speed increases the efficiency of grinding dolomite; no maximum of efficiency appears in the range of speeds tested. The results with chert were inconclusive. If the performance of large mills is indicated

Jan 1, 1951

By Harlowe Hardinge, R. C. Ferguson

Laboratory and plant data covering 12 different operations show that lower than "standard" ball mill speeds increase grinding efficiency. In the case of high pulp-level mills, the gain is so great that the increase in capital cost of the larger lower speed mill will pay for itself in less than a year's time. THE object of this paper is to show the economic advantage of operating ball mills at relatively slow speeds. Although many operators know that grinding costs are reduced by operating mills at low speeds, they seldom do so for a variety of reasons. When capital is limited, first cost is of primary importance, and installation of the less costly, highspeed mill is favored. When materials are in short supply, as during a war period, capacity rather than economy is the primary consideration. Existing mills are then speeded up to provide the desired increase in capacity, and new mills are purchased without consideration of maximum grinding efficiency. When the mill speed is increased to secure more production, the increase in grinding costs is often obscured by the total benefit derived from the additional output. This increase in grinding cost is seldom recognized as being sufficient to be worth correcting. For these reasons one may be lead to the erroneous conclusion that slow-speed operation is only occasionally applicable. The effect of mill speed on grinding efficiency has been studied in laboratory tests."! The results of some of these tests are given in tables I and I1 and fig. 1. In these tabulations, batch tests using ore charges of 125 to 200 lb were considered comparable to high pulp-level operation and charges of 50 to 100 lb comparable to low pulp-level operation. For the high pulp-level operation, these tests indicate the existence of a maximum in efficiency for operation at a speed of 50 pct of critical, both for the soft (dolomite) and hard (chert) ore. In low pulp-level operation, a decrease in speed increases the efficiency of grinding dolomite; no maximum of efficiency appears in the range of speeds tested. The results with chert were inconclusive. If the performance of large mills is indicated

Jan 1, 1951

By R. C. Ferguson, Harlowe Hardinge

Oscar Johnson—In my opinion, the effect of mill speeds on grinding costs must be studied along with capital investment and dollars gathered together as profits. Comparing the entire groups of operators with those who have had the opportunity to make slow-speed mill studies, I think you will find the latter small in numbers. Most managers want the equipment worked to its maximum output. There are, however, some installations where plant and mill sizes are such that they can do the job with reduction of mill barrel speeds. The past and the present installations of the industry are laid out to get the most capacity for the least capital outlay. This is the case even with the plants of Chile Exploration, International Nickel, Morocco, and Anaconda, now under construction or being changed. The industry recognizes that most all equipment it buys today is good and can be depended upon for efficient performance. Under this scheme of things, I am doubtful that slow-speed ball mill operation will be generally applicable. With reference to the U. S. Bureau of Mines laboratory tests, I think table II could have been omitted. It is inconclusive as to maximum efficiency for the low-pulp level mill on hard ore. There should be no question about this point. However, data on mill speeds can be found to substantiate various theories as well as refute them. Gow, Guggenheim, Campbell and Coghill, in their paper on Ball Milling,' believe their 2 x 2 ft laboratory mill reflects results that can be expected from large mills. If so, then referring to their table 11, they state, "The conclusion to be drawn from this second series is that high speed, not exceeding 72 pct of the critical, favors capacity, as before, but that with proper conditions of operation high speeds may give as good efficiency values as low speeds. In this case the efficiency values are nearly constant. A horizontal curve would indicate that the amount of grinding was directly proportional to the power expended, and these tests suggest that such a coildition can be made to exist in commercial operations." Table II (From Paper by Gow et a1)2 Speed. Pot Critical 32 42 52 62 72 82 Capacity: Surface tons per hr (65- mesh) 266 42.1 54.4 65.9 74.3 74.1 Surface tons per hr (200- mesh) 56.1 87.4 112.7 137.1 154.2 153.0 Efficiency: Surface tons per net hp hr (65-mesh) 35.7 36.3 36.3 35.4 34.3 32.3 Surface tons per net hp hr (200-mesh) 75.3 75.3 75.1 73.7 71.0 66.0 Ore in mill, 1.b. 98 100 100 113 122 165 The field performance data, table 111, represents much effort in its collection and preparation. But, one must realize that there are many variables that effect the efficiency of grinding mill operation, and too much must not be assumed as to the effect of some specific change. Possibly with changes in mill speed, the results might be more consistent by also a change in ball rationing, type of ball, volume of ball charge,. p.ulp level and amount of pulp in the mill, pulp consisting, design of liner, circulating load, etc. Also, changes in ore character must be reckoned with when evaluating grinding performance. At present the Climax Molybdenum Corp. is running at much reduced capacity. Mr. James Duggan informs me that at mill speeds of 17 rpm, they save a $0.025 per ton on liners and $0.025 per ton in power, but, if the demand for molybdenum increased, he would go back to higher speed to obtain maximum tonnage, as the values from the increased tonnage would far more than offset the one half saving at the slower speed. The Jnspiration ran a six months' test between mills running 21 rpm and 23.5 rpm. The slower mills ground 10 pct less ore with a slight saving per ton, but when the reduced plant tonnage was checked back into the actual cost figures of concentration, the high-speed mills with their greater tonnage showed considerable advantage. To be convinced of possible practical results from the predictions in the conclusions, I think we would have to rely on the analysis of expert cost accountants to furnish the necessary proof figures. Hardinge and Ferguson are to be commended for the work in preparing this paper. I am convinced that our Massco engineers should go into higher speeds with our equipment. Harlowe Hardinge (authors' reply)—For one, I heartily agree with Mr. Johnson's opening statement that the effect of mill speeds on grinding costs must be studied along with capital investment and dollars gathered together as profits. It was on this basis and for this reason the paper was written. Mr. Johnson, on the other hand, takes the position that, on the whole, low speeds are not justified from the economic standpoint, basing his principal reason on the fact that lower mill speeds cut mill capacities and hence reduce the gross income from the product produced. There is no denying this point. It is almost axiomatic. It is for this very reason that the overall advantage of lower mill speeds has been discounted and even overlooked. It was for this reason mainly that the paper was written in the first place. It is one thing to plan an efficient operation at the outset, basing one's figures on the tonnage requirements at the time, and it is quite another to be confronted with the problem of increasing the output of an existing installation at a minimum of capital expenditure. Economic consideration of a new installation is greatly influenced by referring to an old one. Too often, the analyst assumes that if this practice is followed in the new installation, one would not go wrong. It is just here that he may be wrong. Past practice and low capital expenditure are all too frequently given priority over the engineer's analysis of operating costs. When we are able to start fresh, we should give proper weight to other economic factors which do not exist in an old installation. It is these economic factors that make it possible to spend at the outset just a little more money and get it back in a matter of months and effect big savings for years to come. F. C. Bond—This paper is of considerable importance in that it emphasizes a modern trend to operate ball mills at somewhat slower speeds than formerly. We have checked the data in the paper with that obtained

Jan 1, 1951

By Harlowe Hardinge, R. C. Ferguson

Oscar Johnson—In my opinion, the effect of mill speeds on grinding costs must be studied along with capital investment and dollars gathered together as profits. Comparing the entire groups of operators with those who have had the opportunity to make slow-speed mill studies, I think you will find the latter small in numbers. Most managers want the equipment worked to its maximum output. There are, however, some installations where plant and mill sizes are such that they can do the job with reduction of mill barrel speeds. The past and the present installations of the industry are laid out to get the most capacity for the least capital outlay. This is the case even with the plants of Chile Exploration, International Nickel, Morocco, and Anaconda, now under construction or being changed. The industry recognizes that most all equipment it buys today is good and can be depended upon for efficient performance. Under this scheme of things, I am doubtful that slow-speed ball mill operation will be generally applicable. With reference to the U. S. Bureau of Mines laboratory tests, I think table II could have been omitted. It is inconclusive as to maximum efficiency for the low-pulp level mill on hard ore. There should be no question about this point. However, data on mill speeds can be found to substantiate various theories as well as refute them. Gow, Guggenheim, Campbell and Coghill, in their paper on Ball Milling,' believe their 2 x 2 ft laboratory mill reflects results that can be expected from large mills. If so, then referring to their table 11, they state, "The conclusion to be drawn from this second series is that high speed, not exceeding 72 pct of the critical, favors capacity, as before, but that with proper conditions of operation high speeds may give as good efficiency values as low speeds. In this case the efficiency values are nearly constant. A horizontal curve would indicate that the amount of grinding was directly proportional to the power expended, and these tests suggest that such a coildition can be made to exist in commercial operations." Table II (From Paper by Gow et a1)2 Speed. Pot Critical 32 42 52 62 72 82 Capacity: Surface tons per hr (65- mesh) 266 42.1 54.4 65.9 74.3 74.1 Surface tons per hr (200- mesh) 56.1 87.4 112.7 137.1 154.2 153.0 Efficiency: Surface tons per net hp hr (65-mesh) 35.7 36.3 36.3 35.4 34.3 32.3 Surface tons per net hp hr (200-mesh) 75.3 75.3 75.1 73.7 71.0 66.0 Ore in mill, 1.b. 98 100 100 113 122 165 The field performance data, table 111, represents much effort in its collection and preparation. But, one must realize that there are many variables that effect the efficiency of grinding mill operation, and too much must not be assumed as to the effect of some specific change. Possibly with changes in mill speed, the results might be more consistent by also a change in ball rationing, type of ball, volume of ball charge,. p.ulp level and amount of pulp in the mill, pulp consisting, design of liner, circulating load, etc. Also, changes in ore character must be reckoned with when evaluating grinding performance. At present the Climax Molybdenum Corp. is running at much reduced capacity. Mr. James Duggan informs me that at mill speeds of 17 rpm, they save a $0.025 per ton on liners and $0.025 per ton in power, but, if the demand for molybdenum increased, he would go back to higher speed to obtain maximum tonnage, as the values from the increased tonnage would far more than offset the one half saving at the slower speed. The Jnspiration ran a six months' test between mills running 21 rpm and 23.5 rpm. The slower mills ground 10 pct less ore with a slight saving per ton, but when the reduced plant tonnage was checked back into the actual cost figures of concentration, the high-speed mills with their greater tonnage showed considerable advantage. To be convinced of possible practical results from the predictions in the conclusions, I think we would have to rely on the analysis of expert cost accountants to furnish the necessary proof figures. Hardinge and Ferguson are to be commended for the work in preparing this paper. I am convinced that our Massco engineers should go into higher speeds with our equipment. Harlowe Hardinge (authors' reply)—For one, I heartily agree with Mr. Johnson's opening statement that the effect of mill speeds on grinding costs must be studied along with capital investment and dollars gathered together as profits. It was on this basis and for this reason the paper was written. Mr. Johnson, on the other hand, takes the position that, on the whole, low speeds are not justified from the economic standpoint, basing his principal reason on the fact that lower mill speeds cut mill capacities and hence reduce the gross income from the product produced. There is no denying this point. It is almost axiomatic. It is for this very reason that the overall advantage of lower mill speeds has been discounted and even overlooked. It was for this reason mainly that the paper was written in the first place. It is one thing to plan an efficient operation at the outset, basing one's figures on the tonnage requirements at the time, and it is quite another to be confronted with the problem of increasing the output of an existing installation at a minimum of capital expenditure. Economic consideration of a new installation is greatly influenced by referring to an old one. Too often, the analyst assumes that if this practice is followed in the new installation, one would not go wrong. It is just here that he may be wrong. Past practice and low capital expenditure are all too frequently given priority over the engineer's analysis of operating costs. When we are able to start fresh, we should give proper weight to other economic factors which do not exist in an old installation. It is these economic factors that make it possible to spend at the outset just a little more money and get it back in a matter of months and effect big savings for years to come. F. C. Bond—This paper is of considerable importance in that it emphasizes a modern trend to operate ball mills at somewhat slower speeds than formerly. We have checked the data in the paper with that obtained

Jan 1, 1951

By R. W. Ludt, C. C. DeWitt

The use of froth flotation for the separation of minerals has become one of the most important of ore dressing processes. Its particular adaptability to the enrichment of low grade ores has made the process an important factor in the national economy. The methods have been extended to the recovery of a great number of minerals. Among the few minerals which have resisted efforts toward industrial flotation is chrysocolla, a hydrated partly colloidal copper silicate. Chrysocolla, being a product of natural oxidation, has been found to occur in small quantities with many ores which are recovered by flotation methods. In present practice, these small quantities of copper silicate pass off with the tailings and are lost. The advantages to be gained by a satisfactory process for the recovery of chrysocolla is apparent. Any application of principles which points a way toward the satisfactory industrial flotation process for copper silicate would be of advantage. This paper presents an attack on this problem. Two methods for the recovery of chrysocolla have been developed by the United States Bureau of Mines.1,2 They have been successful on a laboratory scale but have been seriously restricted in industrial application by critical requirements in the procedure. In one of the Bureau of Mines methods,' the ore is activated with sodium or hydrogen sulphide in an aqueous solution at a pH of 4. Amy1 xanthate is then used as a collector with pine oil as a frother in the flotation process. An excess of sulphide acts as a depressant and the state of optimum conditions is difficult to control industrially. In the second Bureau of Mines method,2 soap is used as the collector at a pH of between 8 and 9. The diffi- culties with this process are that soap is not a specific collector, that heavy metal or alkaline earth ions cause the formation of insoluble soaps, and that a more acid solution causes the formation of a free acid which does not act as a collector for chrysocolla. The problem of recovering chrysocolla by flotation involves the selection of a suitable collector. The collector molecule must be composed of an active polar group that has an attraction for chrysocolla, and of a hydrocarbon chain. Certain dyes have been shown to have an attraction for certain minerals. Suida3 found that hydrated silicates are colored by basic dyes. Dittler4 showed that chrysocolla, among other colloidal minerals of acid reaction, preferentially takes up such basic dyes as fuchsin B, methylene blue, and methyl green. Endell5 gave information to show that the colloidal material in clay may be determined by its selective adsorption of fuchsin. A simple experiment, likewise, illustrates the difference in the adsorptive power of chrysocolla and of silica for the basic triphenyl methane dyes. When a mixture of chrysocolla and silica is immersed in a very dilute dye solution, less than 5 ppm, the chryso-colla is rapidly dyed and the silica is dyed more slowly. The difference is substantial but one of degree. Dean2 showed that the dyes, crystal violet and toluidine blue, are taken up by quartz in an adsorption type process. The difference in the adsorptive power, however, offers the means by which a new collector may act. To form such a collector, a hydrocarbon chain must be attached to the dye molecule. This involves a process of organic synthesis. Butyl, hexyl, and octyl hydrocarbon chains were selected for substitution in the malachite green molecule. For the purpose of identification, the alkyl-substituted dyes formed are called: butyl-malachite green; hexyl- malachite green; and octyl-malachite green. An outline of the procedure for their synthesis is given in the appendix. It is generally recognized in the preparation of this type of dye that the chemical structure of some of the dye molecules varies. However, a uniform formula is attributed to the dye. Such a procedure has been followed in specifying the structure of these alkyl-substi-tuted malachite green dyes. The structure is given on the basis of their properties as an homologous series of dyes, on their method of preparation, and on the purity of intermediates used. Structure of substituted alkyl malachite green is: C6H4 N(CH3)2 p-R C6H4 CH C6H4 N(CH)2 Procedure The flotation cell is a Bureau of Mines 100-g, batch unit provided with an air inlet at the bottom above which is a variable speed agitator. The agi-

Jan 1, 1950

By B. V. P. Rao, H. L. Lovell, G. Simkovich

In the study described here, flotation tests were conducted on pure CaF2, on CaF2 doped at several concentration levels with either NaF or YF3, and on natural fluorite samples containing various amounts of impurities in an effort to extend previous knowledge of the influence of crystalline point imperfections on the flotation properties of solids. The results of these tests showed that although recoveries generally decreased as impurities increased, the differences in recovery were not nearly as striking as those detected in earlier experiments. A number of experimentalists have shown an interest in the way in which crystalline point imperfections affect the flotation properties of solids. As long ago as 1953, Welch hypothesized that it might be possible to control surface imperfections and thereby also control the flotation of solids.' A few years later, in some rather extensive work, Plaksin and his coworkers varied the surface imperfections of a number of compounds by oxidation or reduction processes and found that significant variations in flotation properties resulted.2-4 More recently, Simkovich showed that the recovery of both AgCl and PbS could be markedly varied by adding solid-solution impurities, which acted as doping agents, to these compounds.5 Then Mular studied the flotation of ZnO as a function of the doping agents Al2O3 and Li2O.6 Once again, it was found that these impurities, even in very low concentrations, had profound effects upon flotation recoveries. From an imperfection point of view, the defect structures of the compounds tested previously are as follows:7 PbS — amphoteric semiconductor with a Frenkel-type defect structure in the cation sub-lattice.8 AgCl - ionic conductor-Frenkel type defect structure? ZnO — n-type semiconductor, zinc ion interstitials and quasi-free electrons.10 All of these works undoubtedly bear some relation to the findings that the flotation properties of seemingly identical minerals from different ore deposits often exhibit extreme variation in flotation behavior, i.e., that small amounts of analytically undetectable impurities often have a great influence upon flotation

Jan 1, 1969

By Strathmore R. B. Cooke

On the basis of experiments conducted on quartz using a bubble pick-up method, it was shown in an earlier paper1 that this mineral will preferentially adsorb hydrogen, calcium, or sodium ions, depending on the relative concentrations of those ions in the solution in which the quartz is immersed. For quartz particles ranging in size from 0.2 to 1 mm, it was demonstrated that the concentration of calcium in solution (assumed present as ions) necessary to completely activate quartz for flotation is given by the expression: Ca++ = [H+] X 106 + [Na+] X 10-3 In this expression, ionic concentrations are in mols per liter. It was further shown that the pick-up method is apparently more sensitive to changes in reagent concentration than the standard captive-bubble method, and that induction times are apparently much reduced. Since completing these earlier tests, a new type of cell has been constructed; this is shown in Fig 1 and 2. In Fig 1, A is a ground joint, B is a central tube reaching to within a few millimeters of the bottom of the cell, and C is a stopper which may be removed for reagent addition, and serves the further purpose of excluding carbon dioxide from the air during the test. The entire cell is constructed of Pyrex glass, and does not give the trouble experienced with the earlier cell, in which activating ions were released from the glass at high pH values. The only critical factor in the construction of the cell is the clearance between the central tube and the bottom of the cell. This clearance should be sufficiently small that the bubble can be pressed directly on the mineral grains lying on the bottom of the cell. In the pick-up tests to be described in this paper, the reagents used were all of C. P. grade, except the sodium oleate, which was Merck's "neutral powder." The quartz employed was water-clear vein quartz, sized on screens, cleaned with both hydrochloric acid and sodium hydroxide, and given a thorough final washing with distilled water. Experimental procedures were the same as described in the earlier paper. EFFECT OF SIZE OF QUARTZ ON PICK-UP To ascertain the effect of particle size on the adsorption of calcium ions, the quartz was sized from minus 14 plus 20 mesh through the intervening screen sizes to minus 270 mesh plus 400 mesh. Each size was thoroughly cleaned, and then tested in the cell at different calcium chloride and sodium hydroxide concentrations, and at a constant sodium oleate concentration of 20 mg per liter. All particles, within the size range given, exhibited complete pick-up within the curve expressed by the equation above. This presumably means, when the conditions imposed by the equation are satisfied, that this maximum is independent of particle size. However, it was found that the range through which partial particle pick-up occurred progressively broadened as particle size decreased. This is shown in Fig 3, in which curves B, C, and D show the limits at which pick-up just commences (as the pH is increased) for particles of minus 14 plus 28 mesh, minus 65 plus 100 mesh, and minus 270 plus 400 mesh size, respectively. These results indicate that for satisfactory activation, at any given pH, a lower calcium ion concentration is required for fine particles than for coarse particles. EFFECT OF HIGH ALKALINITY ON PICK-UP At calcium concentrations of between 1 and 10 mg per liter, and at high alka-linities, it was noticed that pick-up ceased as soon as calcium hydroxide commenced to precipitate. This effect was investigated at other calcium concentrations, with the same results. Solutions of calcium chloride, containing 105, 104, l03, and l02 mg of calcium per liter were made alkaline with sodium hydroxide until calcium hydroxide just started to precipitate, according to the following equation: CaCl² + NaOH -+ Ca(OH)2 + 2NaCl The beginning of precipitation was taken as that point at which either a faint opalescence appeared in the solution, or a Tyndall cone became ap-

Jan 1, 1950

By K. L. Temple, A. R. Colmer

ACID coal mine drainage presents a peculiarly difficult problem for two principal reasons. First is the fact that the amount of acid water discharged from active and abandoned mines constantly increases as coal fields are developed and thereby creates a steadily worsening nuisance of rapidly growing importance to the general public as well as to the coal industry. Second is the even more significant fact that there is no method yet known that will appreciably alleviate the acid water problem except at totally prohibitive cost to the coal operators. The problem is thus seen to be an urgent one. With a growing need for good quality water and a growing public demand for abatemknt of stream pollution, a large portion of the soft coal industry finds itself in the position of year by year dumping more and more acid into the streams and waterways and having no way in sight of improving the situation or even of keeping it from getting worse. While some studies have been made, it is only in the last few years that the problem has been seriously considered, chiefly by the group of workers under the direction of S. A. Braley at Mellon Institute and by the Bituminous Coal Research, Inc. fellowship at West Virginia University. Hinkle and Koehlerl reported the results of the earlier West Virginia work. This work pointed to certain biological aspects which had not been noticed previously. These may be summarized as two separate findings. One was that sulphur oxidizing bacteria were present in acid mine water and that their possible role in contributing to the acidity should be investigated. The other was

Jan 1, 1952

By G. Gutierrez, D. A. Elgillani, M. C. Fuerstenau

The zero-points-of-charge of apatite, calcite, and fluorite are pH 6.4, 10.8, and 10.0, respectively. Scheelite is negatively charged above at least pH 3. In this article, the flotation responses of these minerals in the presence of potassium oleate and sodium silicate are described and compared with electrokinetic data. Colloidal silica appears to be the species principally responsible for calcite depression, while silicate anion is the species responsible for fluorite depression. Additions as high as 1 x 10-3 mole/liter silicate have no effect on the flotation responses of apatite and scheelite. Selective flotation of nonmetallic minerals is difficult to achieve with fatty acids or soaps by themselves. As a result, specific reagents are added to aid these separations, and one of the reagents commonly employed for this purpose is sodium silicate. Flotation separations of various calcium-bearing minerals such as fluorite from calcite1-3 and scheelite from calcite2,4 and apatite,2,5 for example, almost always involve the use of sodium silicate. The mechanisms by which sodium silicate functions as a depressant are still not understood, probably for a number of reasons. For one thing, the dissolution process of sodium silicate is complex, giving rise to a number of ionic and colloidal species.' Moreover, the type and concentration of these species depend on the ratio of Na2O to SiO2, the concentration of sodium silicate, and the pH of the system.' At the present time, it is not known which species, colloidal silica or silicate anion, is responsible for depression. If colloidal silica is the species that is adsorbing, then adsorption must occur by electrostatic attraction between the colloid and the mineral surface. Silicate anion, on the other hand, may adsorb either physically or chemically. The objective of this paper is to determine first the active species of sodium silicate and then the conditions under which this species will adsorb and function as a depressant. EXPERIMENTAL MATERIALS AND METHODS Pure samples of apatite (Durango), calcite (Iceland-spar), fluorite, and scheelite were used in this investigation. Pure potassium oleate was used as collector, while reagent-grade HC1 and KOH were employed for pH adjustment. The sodium silicates used were samples obtained from the Philadelphia Quartz Co. In one series of experiments, various sodium silicates containing different ratios of SiO2 to Na2O were added to calcite systems. These sodium silicates contained SiO2-to-Na2O ratios of 3.75 to 1, 3.22 to 1, 2.40 to 1, and 1.60 to 1. All other experiments were conducted with the sodium silicate containing the ratio of 3.22 to 1, which is the one that is normally used in industry. Flotation experiments were conducted with 21/2-g charges of 48 x 150-mesh material in conductivity water with an apparatus and technique described previously. Electrokinetic experiments were conducted with both a Zeta Meter and streaming potential apparatus. Particle size was 48 x 65 mesh for the streaming potential experiments. EXPERIMENTAL RESULTS The first series of experiments involved flotation of scheelite in the absence and presence of sodium silicate. As shown in Fig. 1, flotation response was not affected with even the relatively high addition of 1 x 10-3 mole per liter sodium silicate. Interestingly though, no flotation was effected below pH 6 with this collector addition. The responses of apatite to flotation under these same conditions are given in Fig. 2. Similarly, no depression was obtained in basic media under these conditions. Similar experiments were conducted with fluorite, and in this case, depression was noted above about pH 11 with 1 x 10-3 mole per liter sodium silicate (Fig. 3). When calcite was floated with these same levels of addition of sodium silicate, essentially no flotation was possible above pH 7 (Fig. 4). The effect of collector addition with a constant addition of sodium

Jan 1, 1969

By I. J. Lin, A. Mitzmager

This paper describes a series of experiments conducted to determine how the physicochemical properties of the suspending fluid affect the size reduction of quartz in a ball mill. Particular attention was directed toward the differences in rate of grinding and particle size distributions when either polar or non-polar fluids were the suspending liquids. Grinding was performed with water, carbon tetrachloride and methylcyclohexane as the liquids, and also in nitrogen for comparative purposes. From the results ob- tained, the authors have concluded that the presence of water greatly enhances efficiency in the grinding of quartz. Comminution is generally regarded as a purely physical process in which fragmentation of a solid is obtained by the application of suitable forces. In 1944, however, Rehbinder et al.1 pioneered the idea that the interfacial forces between the solid and the media plays an important role in comminution. Since then, a number of researches have shown conclusively that comminution is affected by surface chemical phenomena,2-5 and Gaudin went as far as to compare comminution to a chemical reaction.6 The effect of the viscosity and the density of the fluid media in grinding was also investigated, but

Jan 1, 1969

By F. M. Lewis, E. C. Johnson

Used as a classifier in a plant expanding capacity or changing to closed-circuit operation, the liquid-solid cyclone offers the advantage of being adaptable to existing conditions. This paper presents the costs involved and experience gained from operating a unit of this kind. BY now application of the liquid-solid cyclone to various classification problems in the chemical and metallurgical fields is more or less familiar. A less publicized application receiving wider consideration as time goes on is its use as a classifier in closed-circuit grinding operations. This paper presents the costs involved and the experience gained from operating a unit of this kind. In May 1951 the Tennessee Copper Co. installed in its London mill at Copperhill, Tenn., a liquid-solid cyclone in closed circuit with a 6x12-ft ball mill. The circuit in which this cyclone is used is the regrind or secondary grinding circuit of a copper flotation plant. Originally it consisted of a 7x1-ft conical ball mill in closed circuit with a bowl classifier. Changes made in the primary grinding circuit of the mill necessitated a regrind circuit of greater capacity. The bulk concentrate to be ground was raised from 900 or 1000 tpd to 1200 or 1300 tpd, and to accommodate this increase a 6x12-ft ball mill which had been in the primary grinding circuit was moved into the regrind circuit. When this was done it was found that the classifier was too small to give efficient operation. Space in the plant was at a premium, and since use of the cyclone offered considerable economy the installation shown in Figs. 1-3 was made. The unit installed is a 24-in. diam cyclone classifier with a 20" cone angle. Diameter of the vortex finder is 8 in., the feed entrance area 20 sq. in. Maximum diameter of the adjustable apex valve is 6 in. and minimum diameter approximately 3.5 in. To refer again to Figs. 1 and 2, the cyclone* is mounted motor. It is charged with 18 tons of 1-in. balls. The overflow, or fines, passes by gravity directly to the copper flotation plant. The ball mill discharge and the new feed combine in a sump from which they are sent to the cone by an 8x6 SRL pump at a rate of 497 gpm. Pressure at the cyclone is 5 psi. For reasons dictated by the nature of the ore treated, particularly the bonding of fine grains of chalcopyrite to pyrrhotite, the froth from the second cell of the copper rougher circuit is returned to the regrind mill. This flow, which somewhat complicates the standard closed-circuit calculation, is seen in Fig. 1 to be 156 tpd of solids. This combines with the cyclone classifier underflow to constitute the ball mill feed. The requirement of the regrind product is that it be 77 to 80 pct —200 mesh. The liquid-solids cyclone operating in conjunction with the 6x12-ft mill adequately met this requirement. Of primary importance is the fact that this arrangement provided sufficient classification capacity to replace the undersized classifier and was capable of building up a circulating load to 87 pct, large enough to give an efficient grind. In this instance the circulating load is taken as the ratio of sands to new feed. Any comparison of performance between this liquid-solid cyclone and the replaced classifier would of course be unfair, since the bowl classifier was

Jan 1, 1955

By Lawrence A. Roe

The magnetic reflux classifier, which utilizes the combined effects of magnetic fields and a hindered settling classifier, is a new tool for determining the quantity and quality of middlings in fine-sized magnetite concentrates. Results are given for processing a typical taconite ore, and a sketch of the apparatus is included. IN examining magnetite ores and beneficiated products it often becomes necessary to make critical studies of the amount of grinding necessary to produce the desired degree of magnetite liberation. In the past this has been accomplished by laboratory heavy-liquid tests, which provide a method for selectively removing middling particles and free magnetite from various-sized fractions. Examination of the various products under the microscope results in fairly accurate determination of the degree of liberation. The method is quite efficient on sizes coarser than 325 mesh. Thus the heavy-liquid method of middling separation was satisfactory until the advent of present day magnetic taconite studies. When magnetite concentrates ranging from 70 to 100 pct —325 mesh are studied it becomes apparent that older methods of determining liberation size are not satisfactory and that there is need for a new method. For example, some of the low-grade magnetite ores of the Wisconsin and Michigan iron ranges require grinding to 100 pct —325 mesh to produce a magnetic concentrate containing less than 12 pct silica. Examination of concentrates from such ores often reveals that many of the middling particles consist of only very minor proportions of iron mineral. Thus it becomes important to be able to determine the degree of grinding necessary not only for complete liberation, but also for liberation of only 80, 85, or 90 pct of the total iron mineral content. Actually, complete liberation is never attained, but is often used to designate that degree of liberation necessary for production of high-grade concentrates. A rougher concentrate, produced after elimination of a coarse-sized tailing, can usually be subjected to a second grinding stage and concentrated into a higher grade product than could be produced from the same crude ore with one stage of grinding resulting in the same overall size reduction. This fact adds to the importance of being able to determine partial degrees of liberation on any magnetite ore. Standard laboratory methods such as heavy-liquid separation, microscopic grain counts, Davis tube magnetic separation, magnetic flocculation, classification, flotation, and others often are not applicable, or are prohibitive because of time requirements when large numbers of fine-sized magnetite samples are investigated. The Davis tube magnetic separator is an efficient tool to use in rejecting the non-magnetic mineral particles from an ore sample. The middlings discarded by the tube separator usually are so low in iron content that they can be considered relatively unimportant in liberation studies. This condition is caused by the extremely high flux density used in the Davis tube. This flux density ranges from four to eight times the flux density produced by most of the powerful commercial machines in use today. Thus the problem resolves itself into a search for a method of selectively removing middlings from Davis tube magnetic concentrates which will be both rapid and efficient. Those methods showing most promise in the development of a process for isolating middlings from extremely fine-sized magnetic concentrates were flotation and magnetic flocculation. The use of flotation to remove middlings from magnetic concentrates is reported in the literature.'.' The flotation process is effective in removing middlings from a magnetite concentrate, but physical entrapment of fine-sized free magnetite in the silica-bearing froth is an undesirable feature. The flotation method of removing middlings requires time, effort, and precise control of many variables, and does not meet the required degree of middling isolation. Magnetic Flocculation Magnetic flocculation has long been resorted to"-" in efforts to upgrade magnetite concentrates. One of the new magnetic taconite plants now under construction on the Mesabi Range includes magnetic flocculation in the flowsheet' as an accessory process to remove high-silica middlings and free silica which has been mechanically entrapped in magnetite flocs. The use of magnetic flocculation as a laboratory method of making precise separation of middlings was further investigated, since it offered a rapid, simple method of accomplishing the desired result. Magnetic flocculation involves the subjection of a magnetic concentrate to a strong magnetic field, passing the concentrate in a highly flocculated condition to a hydroseparator or other classifiers of various types, and removing free silica and middlings as overflow products. In an attempt to utilize simple

Jan 1, 1954